Recovery of metal values from complex ores



United States Patent O 3,196,00l RECVERY GE METAL VALUES BRGM CMPLEX@RES @rrin F. Marvin, Cottonwood, Ariz. (1549 VJ. Madison, Phoenix,Ariz.) Filed Sept. 5, 1962, Ser. No. 222,761 7 Claims. (Cl. 75-21) Thisapplication is a continuation-impart of my prior application Serial No.10,210 iiled February 23, 1960, now abandoned.

This invention relates to the recovery of metal values from complexores. It relates particularly to an improvement in the process for therecovery of copper, zinc, lead, iron, silver and gold metal values fromcomplex ores disclosed in my issued Patent No. 2,927,017, dated March 1,1960.

The complex ores, here under discussion, are those that contain two ormore metal values in chemical union or in such physical union as toprevent normal mechanical separation of the values. The complex oresinclude metal values such as copper-iron sulfide, zinc-iron polysuliide,zinc-copper polysullide, zinc-copper-iron trisulfide, lead vanadate,lead chromate, lead molybdate, lead tungstate, lead silicate, and leadcompounds combined with other acidic radicals, and also possibly othercomplex suliides. The ores under discussion also include simplecompounds such as zinc sulfide, copper sulfide, iron suliide, and leadsuiiide uncombined chemically with other suliides, and include alsogangue, dirt and other insolubles.

In the past, attempts have been made to extract copper, zinc and leadvalues from such complex ores by a selective concentration of one of theores and a subsequent smelting to recover the particular value desired.Such a procedure has several major disadvantages. First, the ability toconcentrate a complex ore by standard gravity or oil flotation methodsdoes not meet with much success inasmuch as such processes areadvantageous only in separating the simple suliides from each other(such as copper suliide from zinc suliide), but cannot separate each ofthe metal values in the double suliides, the triple suldes or othercomplex compounds above mentioned. The concentrate, when shipped to thesmelter, thus contains, in addition to the particular metal sought (forexample, zinc), substantial amounts of other values such as copper andlead, and since the smelter is designed to process a particular metal,the shipper does not get paid for these other values (with the exceptionof gold and silver). In fact, if the lead or copper content in a zincconcentrate is above a certain point, penalties for their presence wouldbe exacted by the smelter.

The average assay of so-called zinc and lead selective concentrates forthe Western States in recent years, as produced by standard oil otationmethods, are given in Table l.

Table 1.-A verage assay of selective concentrates It Will be seen bystudying the above table that substantial amounts of copper and leadvalues are present in the zinc concentrates, while substantial amountsof copper and zinc values are present in the lead concentrates. In

igi Faented July 20, 1965 "Ice addition, generally both zinc and leadconcentrates have approximately 10% of iron therein. It can thus be seenthat because of the Vother metal values invariably present, a processwhich involves the steps of rst selectively concentrating an ore for aparticular value, and then smelting to obtain that particular values, isboth inefiicient and uneconomical.

The process of selectively concentrating such complex ores is in itselfa relatively ineiiicient mode of recovery in comparison with that ofmaking a collective concentrate of the values. Thus, it is found that a15-20% increase n the recovery of lead, zinc, copper, gold and silvervalues is possible by collectively concentrating the ore for the variousvalues, instead of attempting to selectively concentrate the ore for aparticular value.

There are still other disadvantages to the smelting operation itself,chiefiy those of the high cost of processing and the high investment perpound of metal recovered. For all of the above reasons, the winning ofcopper, zinc, lead, gold, silver, and iron values from such complex oresas above described, has, in general,4 been commercially unsuccessful.Complex ores treated by the orthodox method outlined above, that is, bysmelting a selective concentrate of each of the metal values, results,on the average, in recoveries of approximately 40-50% copper andapproximately 60-70% zinc.

Attempts at achieving some measure of separation by other means, such asby grinding, have also been unsuccessful since the metal values areusually crystallized in such minute particles that grinding to thedegree of lineness required to actually achieve separation, is verycostly and economically prohibitive.

Wet processes, that is, those involving leaching operations after aroasting operation, have also not been successful in recovering zinc,copper, lead values from such complex ores regardless of whether the oreis first selectively concentrated or collectively concentrated. Byorthodox Wet processes, approximately 60-80% of the total amount of zincand approximately Litl-70% of the total amount of copper present in suchcomplex ores can be recovered.

It appears that the loss in recovery of zinc and copper in both smeltingand Wet processing is due to the formation of certain highly insolublezinc and copper compounds, principally the ferrites of these metalvalues, during the treatment of the ore. In my prior patent, aboveidentified, I disclosed a process whereby substantially completerecovery of copper, lead, zinc, silver, gold and iron values can beobtained from complex ores in a manner that is simple in operation andinexpensive in comparison with both Wet and smelting processes of theprior art. The process described in the above-identified prior patentcomprises the steps of lirst roasting a complex ore at a temperaturesuiciently low to prevent the formation of any substantial amounts ofacid-insoluble compounds of the metal values capable of forming saidacid-insoluble compounds to acid-soluble compounds. Following the irstlow-temperature roasting step, the acid-soluble compounds are separatedfrom the remaining part of the metal values left in the ore residue. Theore residue is then roasted at a temperature suiciently high to convertthe remaining part of metal values such as copper and zinc toacid-soluble compounds.

The process described in my prior patent avoids the formation of highlyinsoluble compounds such as the ferrites of zinc, copper or lead. Thecombination of metal values such as zinc, copper or lead in ferriticcompounds substantially reduces the recovery of these metal values fromcomplex ores. The primary cause of formation of the ferritic compoundsis due to the conversion in an oxidizing roast of ferrous compounds suchas copper by a second leaching step.

encerrar pyrite (CuFeS2) and iron pyrite (FeSZ) into ferrites having aformula MFeZO.,z where M represents a particular metal such as zinc,copper or lead. In the process of my said prior patent, the temperatureof the first roast is maintained at a relatively low value so thatdecomposition and oxidation of the ferrous compounds to ferrie compoundsis attained Without decomposition and oxidation of the other sulfides,either complex or simple, of zinc, lead silver `and gold.V While copperpyrite is decomposed and converted primarily to cupric oxide and somecupric sulfate, no copper ferrites are formed under the conditions ofthe low-temperature roast. It can also .be seen that there is nopossibility of zinc and lead ferrite formation because the compounds inwhich these values rare found are not decomposed and are not, thereforesusceptible to ferrite formation.

While the first roasting step of my prior patented process so convertedthe several compounds in the ore so that the compounds causing formationof ferritic compounds were removed, the first leaching step of suchprocess resulted 'in removal of only about half of the total iron in theore, and the resulting leachate also con'- -tained copper sulfate whichhad to be recovered therefrom. I have now discovered that if the firstlow-temperature oxidizing roast is followed by a reducing roast at asomewhat higher temperature, substantially all of the iron may beleached as ferrous sulfate by means of a vdilute water solution ofsulfuric acid, and that the resulting solution contains substantially nocopper, zinc or lead. It may, therefore, be employed to recover ironvalues, as for example for production of agricultural grade ferroussulfate, without preliminary treatment for separation of copper.

The first low-temperature oxidizing roast of my prior patented processresults in the formation of iron oxide, at best not containing more thanabout fifty percent soluble oxides under operating conditions, aridsince only the ferrous iron is freely soluble in dilute sulfuric acid,the ferrie iron remains behind after the first leaching step. By the useof 4the reducing step in accordance with the present invention, ferriciron is reduced to ferrous iron, and the already present ferrous iron issubstantially unaffected. The reducing step also has the effect ofpartially reducing copper sulfate formed in the first low temperatureoxidizing roast back to metallic copper and the zinc sulfate to zincsulfide, so that they are not removed by the dilute sulfuric acidsolution, While the reducing step may be continued until some elementaliron is found in the roasted ore to assure that reduction has gone tothe point where all of the iron will be removed, in general it ispreferred that no elemental iron be formed during the reducing stepbecause its presence is not conducive to best practice of the inventionfrom a metal recovery standpoint.

The ore residue remaining after the leaching step contains substantiallyall of the identified metal values of interest except y.the iron removedin the leaching step. This residue is then, subjected to ahigh-temperature roast in which the copper and zinc are converted toacid-soluble compounds without danger of ferrite formation.V Theacid-soluble compounds of zinc and copper are separated The residuefollowing the second-leaching step Vis then treated by any of thestandard methods to recover the lead, silver, gold or` otheracidinsoluble rnetal values contained therein.

The advantave of the reducing roast in the process of my presentinvention is that the necessity for treatment of the iirstleachate toremove copper, in particular, is eliminated and the process issimplified by reason of the fact that only .the leachate from the secondleaching step need be treated to recover copper and zinc.

The process of my present invention will become clearly understood byreferring to the following description made in conjunction with theaccompanying drawing which is a diagrammatic flow sheet of the processaccording to my present invention.

ln general, .after the collective or bull: concentration of the metalvalues, the process comprises a first lowtemperature roasting stepwhereby some or all of the copper values in the bulk concentrate arerendered acidsoluble and undesirable acid-insoluble compounds of themetal values are prevented from forming. The roasted ore is thensubjected to a reducing roast, and the ore from the reducing roast isleached to remove the soluble iron compounds. However, apart from iron,the other metal values of the complex ore remain in the leached residue.

The leached residue is then roasted at a substantially highertemperature than the first roast, inasmuch as the undesirable substanceswhich would formV acid-insoluble compounds such as ferrites have beenremoved by the rst leaching step. The zinc and copper are converted intothe acid-soluble form by the second roasting step and leached out by anacid solution.

The lead is then recovered from the remaining residue by suitableconversions and leaching operations, as will be described. After theremoval of the lead, copper and zinc values, the remaining residue istreated to recover the gold and silver by any standar-d method, such asby cyaniding to dissolve the silver and gold therefrom. Any remainingiron values are then removed from the remaining residue by standardreduction of ferric oxide contained therein.

It will be understood that the process of my present invention can alsobe used for the recovery of cadmium, antimony, bismuth, arsenic, cobaltand nickel from complex ores in addition to the metal values set forthabove. The steps for recovery of Cd, Sb, Bi, As, Co and Ni values, whilenot herein described, will be apparent to thoses skilled in the art fromthe description of the process of my invention which follows.

Referring now to the flow sheet, complex ore as mined is iirst groundand sent to a bull: concentrating zone lil along the line l2 where theore is concentrated by standard oil flotation methods. That is to say,la concentrate of the complex zinc, copper, lead iron, and preciousmetal suliides, as well as all the simple sul-des of the metal, isseparated from the gangue, rock and dirt, designated generally by thetailings 14. It is found that the bulk concentrate produced containsapproximately 15-20% more of the metal values than are present in theconibined totals of selective concentrates for each of the values. Thereason for the increased recovery of the zinc, lead, or copper-zincvalues is that standard oil flotation methods for selectivelyconcentrating each of these values separate only the simple suliides ofeach of the elements, and, of course, cannot separate or decompose thecomplex disuliides and trisulfides which are'lpresent'in substantialamounts. It is also possible to employ my process Without any oreconcentration, i.e., processing the ore directly as mined, but thecollective concentration is found much more feasible.

The bulk concentrates, upon analysis, have been found to usually containthe following compounds: Copper-zinc polysuliide Copper-zinc-ironpolysuliide Copper-iron sulfide Zinc-iron polysuliide Copper sulfide YSilver sulfide Zinc sulfide Lead sulfide Lead vanadate Lead chromateVLead molybdate Lead tungstate Lead silicate Lead plus other `acidicradicals l Ferrous suliides, iron disuliides or sulfides or iro Cadmiumsulfide Antimony sulfide t Arsenic sulfide i:

eyery .small amounts por traces,

Bismuth sulfide Gold (elemental)* *Very small amounts o1' traces. Theproportions of the metal values in the bulk concentrate, of course, varya great deal. By way of illustration, however, an average assay of abulk lead concentrate of an Arizona ore is given below.

Metal Percent Zinc (Zn) 9.95 Copper (Cu) 1.16 Lead (Pb) 28.8 Iron (Fe)21.0 Silver (Ag) (OL/ton) 38.3 Gold (Au) (oz./ton) 1.0 Cadmium (Cd) 0.01Antimony (Sb) 0.1 Silicon (Si) 2.0

The bulk concentrate is sent along a line 16 to a first low-temperatureroasting zone 18, the temperature of which is maintained between certainspecified, relatively low limits. The primary purpose of thislow-temperature roast is to prevent formation of ferrites which wouldotherwise combine with the zinc, copper and lead values in the oreconcentrate to form acid-insoluble ferritic compounds. The term roast,roasting and the like used herein and in the claims, are used to denotea heating in an oxidizing atmosphere.

Preferably, the temperature in roasting zone 18 is maintained above 340C., but below 460 C. While the formation of ferrites is minimized in thetemperature range from 340 C. to 400 C., no substantial ferriteformation occurs when the temperature of the iirst roast is in the rangefrom 340 C. to 460 C. The upper portion of this latter range has theadded advantage that the optimum temperature of the first roast formaximum solubilization of iron in acid appears to be around 454 C. Whenthe temperature of the first roast is significantly increased above thistemperature, the acid-soluble iron in the roasted concentrate decreases.For example, at a temperature of about 650 C. only about 10% acidsolubleiron is produced, whereas, at 454 C. from 50% to 80% of the iron presentis soluble in 15% sulfuric acid solution. Similarly, when thetemperature of the roast is decreased below about 454 C., the percentageof acid-soluble iron in the concentrate is decreased. For example, atabout 370 C. approximately 25% of the iron present in the concentrate isacid-soluble. So long as the formation of ferrites is avoided, it ispreferable in the practice of the process to take advantage of a iirstroasting temperature at which maximum solubilization of iron results.

At the low temperature of roasting zone 18, while no copper ferrites areformed, copper pyrite is converted to cupric oxide and, to some extent,cupric sulfate. In addition, the ferrous compounds are readily convertedto acid-soluble iron compounds and to oxides at these temperatures.

Zinc sulfide, as well as the sultides of cadmium, bismuth, cobalt,nickel and silver remain substantially unchanged at the low roastingtemperature of roasting zone 1S. Lead sulfide is partially converted tothe sulfate, but the major portion of this sulfide remains unchanged.The compounds of lead with an acidic radical, such as lead vanadate,lead chromate, lead molybdate, lead tungstate, etc., are not convertedto the oxide or sulfate form, to any appreciable extent, during thelow-temperature roast and remain substantially unchanged.

The decomposition of the simple and complex sulides of iron and copper(principally copper pyrite and pyrite) during the roasting oeration inzone 1S produces sulfur dioxide, which is used as the basis for themaking of sulfuric acid for the acid leaching operation to be described.'I'he sulfur dioxide gas leaves the low-temperature roasting zone alonga line 20.

The roasted concentration is sent to a reducing zone 22 along a line 24.To provide a reducing atmosphere, a reducing medium such as carbonmonoxide (CO) is added to the reducing zone through a line 26. Thecompounds of iron in the roasted concentrate are reduced in accordancewith the reactions previously set forth. Ferrie oxide (Fe203) is reducedto ferrosoferric oxide (Fe304) which in turn is reduced to ferrous oxide(FeO). The ferrous oxide is reduced to elemental iron. In general,ferrie oxide and ferrosoferric oxide are reduced to ferrous oxide beforeany appreciable amount of ferous oxide is reduced to elemental iron.

I have already pointed out that while the reducing roast may becontinued until some elemental iron is formed to assure maximumreduction of ferrie iron to ferrous iron, the ideal condition is one inwhich all of the iron is reduced to ferrous iron and no metallic ironhas been formed. Such a condition would, of course, permit removal ofsubstantially one hundred percent of the iron during the first leachingstep. The degree of completion of the reducing roast, that is, theextent to which all of the iron is reduced to a condition for leachingby dilute sulfuric acid, is generally controlled by the variables oftime, temperature, pressure and concentration of reactants and products.Those skilled in the art are familiar with the manner in which theseconditions may be controlled to yield a given desired result. Ingeneral, a reducing roast in the range of 450 C. to 650 C. in anatmosphere of hydrogen or -other common reducing material commonly usedin metal roasting operations has been found satisfactory. The removal ofwater and carbon dioxide from the reducing zone through line 28 removesthe products of the reduction reaction and promotes the rate thereof.

The concentrate from reducing zone 22 is delivered to a leaching zone 30along line 32, and leached with sulfuric acid of from 1% to 50%concentration. The sul furic acid leach solution is delivered to zone 30by line 34, and is preferably maintained at a slightly elevatedtemperature, for example, at 60 C. This sulfuric acid leach dissolvesall of the ferrous iron in the ore afected by the previous lowtemperature oxidizing and higher temperature reducing roasts. Ingeneral, this usually comprises almost all of the iron in the ore.Substantially no copper or zinc are removed from the ore by these steps.The resulting leach solution containing principally ferrous sulfateleaves the leaching zone by line 36. Since the first acid leach solutioncontains principally iron as ferrous sulfate, it may be discardedwithout recovery of the iron if the cormnercial situation does notjustify treatment to recover the iron. 1f, however, recovery of the ironvalues is advantageous, the solution is passed to an Iron removal ZoneA, indicated by reference character 33, in whichl the ferrous iron isoxidized by passing an oxygen-containing gas, such as air, through thesolution. The oxygencontaining gas enters Iron Removal Zone A throughaline 40. Approximately half of the iron is precipitated as ferricsulfate [Fe2(SO4)3] which leaves Iron Removal Zone A along a line 42.

The remaining ferrie iron passes to an Iron Recovery Zone B, designatedby the reference character 44, along a line 46, Where it is precipitatedby the addition of zinc oxide as hydrated ferrie oxide (FegOB-HZO).Also, traces of any other base metals that may be present are therebyprecipitated, with the exception of possible traces of magnesium,manganese and sodium. The zinc oxide employed in precipitating theferric iron is obtained by later steps in the process, as will bedescribed, either during the formation of the zinc metal or directlyfrom the calcines formed in the second roasting step.

The final solution remaining comprises essentially zinc sulfate only,and passes from Iron Recovery Zone B along a line 48 to a Zinc RecoveryZone to be described.

The concentrate not dissolved by the iirst leaching solution is sent toa second high-temperature roasting zone 7 50 along a line 52 and isroasted at a temperature above 540 C., preferably within the range of600`to 800 C. A minimum temperature of 540 C. is required since, Whilethe minimum decomposition temperature of the lead sulfide is 450 C., theminimum decomposition temperature of the zinc sulfide is 540 C. In thesecond high-temperature roasting step, the zinc sulfides are convertedto the oxides, with some being converted to sulfates; the copper isconverted to copper oxide, with some beingy converted to sulfatos; theremainder of the complex sulfides are decomposed and oxidized; likewise,cadmium, antimony, bismuth, arsenic, cobalt, and nickel sulfides and thelike are similarly converted. The lead sulfide present is convertedtoboth the oxide and the sulfate. Silver sulfide (AgS) is readilydecomposed and reduced to the elemental form. lt should be noted thatsince all ferrite forming components have been removed by the firstleach, the high-temperature roast can be conducted without fear of Vlossof any zinc, copper, lead, iron or other metal values such as Cd, Sb,Bi, As, Co and Ni as insoluble ferrites.

Sulfur dioxide is also produced in the second roasting zone and ispassed from zone Sti through a line 54 to an acid plant for use inmakingsulfuric acid. The sulfurie acid is then used in first leachingzone 30 or in a second leaching Zone now to be described.

n The concentrate from the second roasting zone 50 is sent along a lineo to a second acid leaching zone 58 where a sulfuric acid solution offairly high concentration, e.g., between l550%, is introduced through aline 60 to leach substantially all of the zinc and copper values in theconcentrate. ATraces of ferriev iron values may go into the leachsolution at this point. The leaching operation may be conducted in asingle or multiple number of stages, as the particular ore concentratemay require.

inasmuch as the concentrate after the second roasting, depending on theconditions used for the roast, may contain essentially all of the zincas Zinc oxide, an alternative procedure sometimes followed is to taketherequisite portion of the concentrate directly to Iron Recovery Zone B atdi. for the removal of iron and metal impurities by substitution of theiron and base metals for the Zinc in zinc oxide. Whether this mode ofpurification of the first leach solution and recovery of the metalvalues therefrom is chosen, or whether zinc oxide produced from theoxidation of the zinc pigs during production of zinc in later stages ofmy process is used, is an economic question.

in following this alternate process Vand employing the calcines from thesecond roasting zone, a portion thereof is sent along dotted line 62 toIron Recovery Zone B at 44. The zinc-iron exchange takes place in IronRecovery Zone B and the resulting zinc sulfate solution leaves alongline 48.

The second acid leach solution is separated from the solid concentratesand leaves zone 5S along a line 64, the second acid leach solutioncontaining all of the zinc and copper values and all the cadmium,bismuth, cobalt and nickel (present in trace amounts). The copper,cadmium, bismuth, cobalt and nickel values are removed from the secondleach solution by any appropriate means, such as by oxidation-reductionreaction with zinc dust entering a Copper Recovery Zone 65 along a line68. The copper recovered is sent, along with the traces of otherelements aforementioned, to refining operations through a line 70.

The remaining solution is comprised substantially of zinc sulfate fromtwo sources, zinc sulfate produced by the second leaching step in zone58 and the Zinc sulfate produced by oxidation of added zinc metal in theCopper Recovery Zone at66. The solution is sent along a line 72 directlyto a Zinc Recovery Zone '74 or by-passed to Iron Recovery Zones A and Bat 38 and 44, respectively, if the presence of iron and other impuritiesso requires. That is to say, the Zinc sulfate solution may containtraces of ferrous (and ferrie) iron which can be removed @.9 byoxidation of the ferrous to the ferrie formin Iron Recovery Zone A at3S, and by neutralizing with zinc oxide in the iron Recovery Zone B at44, to thereby precipitate all the iron as ferrie sulfate and hydratedferrie oxide (and to precipitate other impurities, if present), asdescribed previously. Thus, if the concentration of ferrous and -ferriciron or other impurities is sufficiently high to possibly interfere withsubsequent zinc recovery (a maximum of about 0.007%), the zinc sulfatesolution is sent to the aforementioned Iron Recovery Zones A and B alonga dotted line '76.

T he zinc sulfate solution, leaving Iron Recovery Zone B along line d8and/ or leaving Copper Recovery Zone 66, then enters the main zincsulfate line "i8 to be sent to Zinc Recovery Zone 74. The zinc isrecovered by any of a variety of methods, for example, byelectro-deposition. `The spent electrolyte, containing some zinc sulfateand sulfuric acid, is then recycled along aline to second leaching zone53 for use in the leaching operations therein, as previously described.V

The presence of magnesium, sodium calcium and potassium is usually sosmall as not to interfere with the electro-deposition. This isespecially true if the first leach solution is not processed for thesmall amount of Zinc it may contain. lf, after some time, however,substantial buildup of these elements doesoccur (they can be removedprior to deposition of the zinc by any of a number of standard methods.The presence of manganese in the electrolyte is preferable inasmuch asit will act as a built-in7 depolarizino agent during zincelectrodeposition.

If one were to follow the prior art practices and conduct a single,high-temperature roast, the average recovery of copper is found to be inthe neighborhood of t0-70%, whereas by following the two-stage roasting,reducing and leaching operation described above, recovery of copper isalmost always 99+%. The high recoveries are thought to be due entirelyto the prevention of the formation of copper ferrites which areacid-insoluble compounds and the chief source of loss in orthodoxleaching and smelting operations.

Similarly, with respect to zinc, by following the othodox one-stageroasting operation at high temperatures, the average recovery of zinc,regardless of whether it is smelted from a selected concentrate orWhether it is leached, is in the neighborhood of oil-80%. The loss isdue primarily to the formation of ferrites or other undesirablecompounds formed during the roasting operation. The staging of thelow-temperature roast and the high-temperature roast with theintermediate removal Vof ferrous iron, eliminates the undesirableformation of zinc ferrites and substantially increases the percentage ofrecovery of zinc so that it is substantially a complete recovery.

The Zinc recovered in Zinc Recovery Zone 74- is sent along a line 82where it is melted into pigs for shipment. A portion of the zinc isground into dust to be sent through a line'St, to Copper Recovery Zone66 for the oxidationreduction reaction therein.

`In melting the zinc to pig, there will be some oxidation of the Zinc tozinc oxide (ZnO). The Zinc oxide thus produced is sent from line S2 toIron Recovery Zone B at 44 through a line do for the precipitation ofimpurities and iron as hydrated iron oxide, as previously explained.

The leached concentrates containing all of the gold, silver, lead, andthe majority of the iron values insoluble in the first leach leave thesecond leaching zone along a line Sg and enter a Lead Conversion Zone90.The lead values are substantially entirely present in the form of leadsulfate because of the prior sulfuric acid leaching steps in the firstand second leaching zones. In order to leach the lead values from thesolids and s-tillv regenerate the leaching agent, the lead is firstconverted 4to the carbonate form in Lead Conversion Zone andisthereafter leached out as lead acetate with acetic acid. The leadacetate is then converted to sulfate by means of sulfuric acid wherebythe lead sulfate is recovered as a precipitate and the acetic acid isregenerated for use in leaching the lead sulfate from the solidconcentrates.

To this end, a near-boiling sodium carbonate solution enter-s zone 90along a line 92 at -a concentration of 10 to 40% and converts the leadsulfate in `the solids to insoluble lead carbonate. A sodium sulfatesolution thus leaves zone 90 along a line 94 .according to the equation:

The lead-containing solids are then sent from zone 90 yalong a line 96to a third leaching zone 9S wherein the solids are leached with anacetic acid solution entering zone 98 along a line 100. The leadcarbonate is thus decomposed and soluble lead acetate solution is formedaccording to the equation:

The lead then leaves the third leaching zone along a line 102 lto a LeadRecovery Zone 104. Snlfuric acid is then added Ithrough a line 106 tothe solution of lead acetate in zone 104 whereupon the lead precipitatesas lead sulfate, leaving the recovery Zone along a line :108.

The -ace-tic acid solution is regenerated according to Athe equation:

leaving zone 104 along a line 110 to be recirculated to third leachingzone 98. Fresh acetic acid solution is added to the recirculating aceticacid line D-110 along a line 112.

It should be noted that lead may be recovered from the system in otherforms lthan the sulfate, inasmuch as the addition of any mineral acidwhich is more highly ionized than lead acetate will precipitate the leadas the lead combined with acidic radical of the miner-al acid employed.It :should be noted valso that lead cau Ibe recovered as a sponge metalfrom the Lead Recovery Zone by electrolysis, .if it is more desirable tosell the lead in this form.

The solids, Aafter undergoing the acetic acid leaching treatment in thethird leaching zone at 98 are sent along a line 114 to a fourth leachingzone 116 for the purpose of leaching out the silver and gold Values fromthe remaining insolubles in the solids (such as ferrie oxide, andsilica) by-sEudard rneans. For example, sodium cyanide enters leachingzone 116 along aline 118 leaching out the silver and gold values assilver cyanide and gold cyanide, respectively. The soluble cyanidevalues leave the fourth leaching zone along aline 120 to be sent to aGold and Silver Recovery Zone `122. It is to be emphasized that the goldand silver entering -the fourth leaching zone are pure gold and silvermetal, respectively, which are readily converted to the cyanide complexform. Thus, a short leaching time only is required. The gold and silvervalues are selectively converted to 1.their reduced elemental form byreduction with zinc rnetal or by other standard methods.

The iron remaining in the concentrate, which is in the form of ferrieoxide, as Well yas other insolubles such as silica are sent along a line124 to appropriate reduction processes where Ithe iron is recovered inan Iron Recovery Zone C 126. For example, the iron oxide may be reducedin an hydrogen atmosphere or by carbon monoxide gas in a blast furnace.The silica-tes and other remaining insolubles are removed from vironRecovery Zone C along a line 12S.

An example which is typical of the process above-described is set forthbelow.

10 EXAMPLE A ground sulde ore was colle-ctively concentrated with theraw concentrate having the yfollowing assay:

Metal value: Percent by weight Zinc 39.0 Copper 1.0 Lead 12.0 Iron 12.0Cadmium 0.1

From the above assay, it will be seen that the concentrate was high inzinc and iron.

The concentrate was roasted in a low-temperature roasting zone in anoxidizing atmosphere containing about 8% sulfur dioxide. The roastingtime was about two and one-:half (2l/2) hours and the temperature of theroast was approximately 400 C.

The roasted concentrate was then subjected to a reduc- .ing roast wherehydrogen gas provided the reducing rnedium. The residence Atime in thereducing roast was about 0.25 hours and the tempera-ture Aof thereducing roast was about 538 C.

rthe concentra-te from the reducing zone was then leached with 4a 15sulfuric acid solution at a temperature of abou-t 77 C. for one hour.The leach solution was separated from the ore residue and ,the leachsolution was Itreated to recover the metal values contained therein. Therecovery of metal values from the rst acid leach solution is summarizedin the table below:

Percent by weight extracted Mea mercenaire Zinc 1.0 Copper 0.5 Lead 0.0Iron 95.04- Cadmiuin 0.0

in second leach of total Metal VallJESI in Ore residue Zinc 990+ Copper99.0-1- Lead 0.0

Iron 50.0 `Cadmium 99.04-

The ore residue was then 'treated for recovery of lead with a 15%solution of sodium carbonate at 100" C. and a 10% acetic acid solution,as previously described in conjunction with the tlow sheet. The Aresiduefrom the lead recovery zone was not further treated to recover silver,gold or .other residual metal values.

The tot-al recovery of metal values from the ore concentrate issummarized in the table below:

Percent total recovery based on Weight in raw .f Metal values. oreconcentrate Zinc 98.0-1- Copper 95.0%- Lead 98.0 Iron 97.5 Cadmium98.0-1-

It will be seen from the above example that the recovery of metal valuesis substantially complete based upon the metal values initially in theraw concentrate. Furtherclaims:

more, it will be seen that only iron values are found in the first acidleach solution and that substantially all recovery of copper and zincvalues is from the second acid leach solution. Where, therefore,recovery of iron is not sought, the process may be economicallypractised by treating only the second acid leach solution for recoveryof metal values. Y

As the specification makes clear, sulfide ores of the characteridentified as a'rule are mined for their zinc, copper and/ or leadvalues and normally iron, while usually present in substantialproportions, is not considered of prime importance in commonly usedrecovery procedures. Gold and silver, when present in economicallysignificant proportions, are also usually recovered. Such elements asvanadium, chromium, molybdenum, tungsten, cadmium, arsenic, bismuth andselenium are frequently present in relatively small proportions asimpurities, but frequently not in sufficient quantity to warrantrecovery. The process of the present invention, however, does permitrecovery of some ofrthese materials profitably even when present in irelatively small amounts. y

Arsenic, antimony and selenium, for example, when present in the orewill normally be leached and recovered as sulfates during the firstleaching step. Removal of antimony at this stage may be importantbecause among the acid insoluble compounds which may be produced arelead antimoniate, wh-ich could be formed by the final oxidizing roast ifantimony were not previously removed.

The present invention -is particularly useful in the treatment ofcomplex sulfide ores of the type identified not only from its ability toproduce separate concentrates of such values as copper, zinc, lead andthe like, but in the removal of substantially all of the iron and otherelements such as antimony capable of forming insoluble compounds withcopper, zinc, lead and the like, so that such metal values may later berecovered either 'by the process of the present'invention or optionallyby known processes of thev prior art.

While a preferred embodiment of this invent-ion has been described, itwill be understood by those skilled in the art that changes andmodifications may be made that lie within the scope of the invention asdefined by the I cla-im:

l. A process for recovery of metal values from a cornplex sulphide oreof the character identified and including copper, lead, zinc and iron,which comprises:

(A) roasting said complex ore at a temperature sufficiently low toprevent any substantial formation of ferrites but sufiiciently high toconvert at least part of i2 the iron values to ferrie and ferrousoxides, and leaving other metal sulphides unoxidized,

(B) heating said complex ore in a reducing atmosphere to convert thesaid ferrie oxide to ferrous oxide (F60),

(C) leaching said complex ore to separate therefrom the ferrous oxide.

2. The process claimed in claim 1 wherein the leached residue is roastedat a temperature sufficiently high to convert at least a part of theremaining metal values to acid soluble compounds.

3. A process for recovery of metal values from a cornplex sulphide oreof the character identified and including copper, lead, zinc and iron,which comprises:

(A) roasting said complex ore at a temperature below 400 C. to convertat least a part of the iron values to ferrie and ferrous oxides, andleaving other metal sulphides unoxidized,

(B) heating said complex ore in a reducing atmosphere above 550 C. toconvert said ferrie oxide to ferrous oxide.

4. The process claimed in claim 3, wherein the said ferrous oxide isacid leached to separate the iron from the other constituents.

5. The process claimed in claim 4, wherein the acid leached residue isroasted at a temperature sufficiently high to convert the remainingmetal values -to loxide compounds.

6. The process claimed in claim 5, wherein the oxide compounds are acidleached to recover the copper and zinc and leaving an insoluble residuecontaining all of the lead.

7. The process claimed in'clairn 6, wherein the concentrate remainingafter the sec-ond acid leach is treated with sodium carbonate solutionto thereby convert the lead in said concentrate to lead carbonate, theresulting lead carbonate is then treated with acetic Vacid to leach thelead values from the concentrate as lead acetate, and re-precipitated bytreatment of said lead acetate with a mineral acid whereby said aceticacid is regenerated for reuse in the leaching of lead carbonates fromsaid concentrate.

References Cited by the Examiner UNITED STATES PATENTS 153,573 7/74Kidwell 75-21 1,637,838 S/27 Simonds 75-21 1,833,686 ll/31 Meyer 75-12,123,240 7/38 Hammarberg 75-21 2,927,017 3/60 Marvin 75-101 3,053,6519/62 McGauley 75-1 BENIAMIN HENKIN, Primary Examiner.

1. A PROCESS FOR RECOVERY OF METAL VALUES FROM A COMPLEX SULPHIDE ORE OFTHE CHARACTER IDENTIFIED AND INCLUDING COPPER, LEAD, ZINX AND IRON,WHICH COMPRISES: (A) ROASTING SAID COMPLEX ORE AT A TEMPERATURESUFFICIENTLY LOW TO PREVENT ANY SUBSTANTIAL FORMATION OF FERRITES BUTSUFFICIENTLY HIGH TO CONVERT AT LEAST PART OF THE IRON VALUES TO FERRICAND FERROUS OXIDE, AND LEAVING OTHER METAL SULPHIDES UNOXIDIZED, (B)HEATING SAID COMPLEX ORE IN A REDUCING ATMOSPHERE TO CONVERT THE SAIDFERRIC OXIDE TO FERROUS OXIDE (FEO), (C) LEACHING SAID COMPLEX ORE TOSEPARATE THEREFROM THE FERROUS OXIDE.